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Article

Gold Leaching from an Auriferous Ore by Alkaline Thiosulfate–Glycine–Copper Solution

by
Alex S. Redrovan
*,
Ernesto de la Torre
and
Carlos F. Aragón-Tobar
Department of Extractive Metallurgy, Escuela Politécnica Nacional, Ladrón de Guevara E11-253, P.O. Box 17-01-2759, Quito 170525, Ecuador
*
Author to whom correspondence should be addressed.
Metals 2025, 15(2), 204; https://doi.org/10.3390/met15020204
Submission received: 17 December 2024 / Revised: 10 January 2025 / Accepted: 13 January 2025 / Published: 14 February 2025
(This article belongs to the Special Issue Advances in Mineral Processing and Hydrometallurgy—3rd Edition)
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Abstract

:
The thiosulfate–glycine–copper system has emerged as a promising alternative for gold recovery, offering significant advantages over cyanidation and ammoniacal thiosulfate leaching. Recognizing the limitations of thiosulfate degradation in ammoniacal systems, this study focused on optimizing the thiosulfate–glycine–copper system for gold recovery using an auriferous ore with (10 g t−1) of Au. The ore was associated with aluminosilicates such as grossular (64%) and clinochlore (12%). Leaching conditions were systematically varied, including thiosulfate (0.5–1 M), glycine (0.3–1.75 M), copper sulfate (2–10 mM), pH (9.3–10.5), temperature (20–60 °C), 6 h, and potassium permanganate concentrations (0.004–0.04 M), and dosing intervals were also optimized. Thus, the best conditions were thiosulfate (0.7 M), glycine (1.75 M), copper sulfate (5 mM), pH 9.3, 60 °C, and permanganate addition every 2 h. This system achieved 89.3% gold recovery in just 6 h, comparable to cyanidation (89.8% in 24 h) and ammoniacal thiosulfate (58% in 6 h), but without generating toxic effluents, such as in the cyanidation process. Additionally, a gold dissolution mechanism was proposed, highlighting glycine’s role in stabilizing cupric ions and enhancing thiosulfate efficiency. This study underscores the thiosulfate–glycine–copper system as a sustainable and effective method for gold recovery.

Graphical Abstract

1. Introduction

1.1. Traditional Methods for Gold Recovery

The most widely used industrial method for the recovery of gold from auriferous ores is sodium cyanide leaching, due to its low reagent and high recovery (89%) [1,2]. However, its use poses serious environmental risks due to the high toxicity of the generated effluents, which can lead to severe ecological and public health consequences [3,4,5]. For this reason, alternatives, such as sodium thiosulfate, have been extensively studied for their lower cost and reduced toxicity [3,6].
Despite its advantages, the industrial scale use of thiosulfate presents two main challenges: first, the degradation of sodium thiosulfate into by-products such as tetrathionates, polythionates, and sulfates, which increases reagent consumption [2,7]; and second, the low affinity of the gold–thiosulfate complex to adsorb onto activated carbon [6,8]. To mitigate the first problem, a catalyst, such as stable cupric ions, is typically used to accelerate the dissolution rate of gold [5,9,10]. However, cupric ions can be reduced to cuprous ions by dissolved oxygen in the solution, which in turn accelerates the degradation and consumption of thiosulfate [11,12]. To prevent this, cupric ions are generally stabilized using ammonia, which forms the “cuprotetramine” complex, keeping copper in its Cu (II) state, thereby facilitating the formation of the gold–thiosulfate complex [6,9]. Although ammonia improves gold recovery, its use in high concentrations poses environmental risks similar to cyanide [4,13].
On the other hand, the recovery of dissolved gold from thiosulfate solutions using activated carbon presents significant limitations [14,15,16]. The low affinity of the gold–thiosulfate complex, due to the difference in polarity with the carbon and the molecule’s size relative to the pores of conventional activated carbon, hinders this process. In contrast, the traditional method using activated carbon achieves up to 98% adsorption of the gold–cyanide complex [15,17,18,19]. However, it has been demonstrated that modifying the specific surface characteristics of activated carbon, by introducing nitrogen-containing functional groups, can significantly improve the gold adsorption rate [14,20].

1.2. Thiosulfate–Glycine System: An Alternative for Gold Extraction

As mentioned in the previous Section, gold leaching with sodium thiosulfate solutions requires stable cupric ions to catalyze the dissolution of gold. Given the toxicity of ammonia, it is necessary to find an effective substitute [21].
Glycine, a non-toxic and low-cost amino acid, is commonly used in the selective recovery of copper from oxidized ores. It is primarily used as a dietary supplement or a regulator of the central nervous system [22,23]. However, [22,24,25] demonstrated that selective leaching from pure gold sheets and copper–gold concentrates, using glycine in synergy with other amino acids or cyanide, achieved 70% gold recovery within 50 h. This recovery is due to glycine’s ability to form strong complexes with Cu2+, Co2+, and Ni2+ [26,27], thereby reducing reagent consumption. Particularly with Cu2+, the copper–glycinate complex is more stable than the ammoniacal copper complex “cuprotetramine” [4,25,28]. Furthermore, [1,29] demonstrated glycine’s ability to dissolve gold, recovering 50% of the gold and silver under alkaline conditions and in synergy with oxidizing agents, such as hydrogen peroxide or potassium permanganate. However, glycine leaching conditions—temperature (60 °C), pH (11), and leaching time (167 h)—reveal that the leaching rate is very slow [1,23].
For these reasons, the use of glycine within the sodium thiosulfate system emerges as a novel and underexplored alternative for the metallurgical treatment of gold ores. It can replace ammonia, reducing thiosulfate consumption; additionally, it acts as a co-leaching agent for gold [5,27,30,31,32].
The objective of this research is to develop an alternative leaching system using sodium thiosulfate. For this purpose, an Ecuadorian gold-bearing ore was used to optimize the thiosulfate–glycine–copper system by varying the concentrations of sodium thiosulfate, glycine, copper sulfate, potassium permanganate, pH, and temperature, until achieving recoveries comparable to those obtained with cyanidation and ammoniacal thiosulfate leaching.

2. Materials and Methods

2.1. Materials

The chemicals used in this study were analytical grade copper sulfate pentahydrate (AppliChem Panreac, 99%, Darmstadt, Germany), anhydrous sodium thiosulfate (LOBA Chemie, 98%, Mumbai, India), glycine (Sigma-Aldrich, 99%, St. Louis, MO, USA), sodium hydroxide (Merck, 99%, Darmstadt, Germany), and technical grade potassium permanganate. All experiments were conducted with distilled water, and pH control was achieved using a 20% v v−1 ammonium solution for the ammoniacal thiosulfate system and a 4 M sodium hydroxide solution for the thiosulfate–glycine–copper system.

2.2. Methods

2.2.1. Ore Characterization

The auriferous ore used in this study was obtained from the Guayzimi mining area, located in the province of Zamora Chinchipe, southern Ecuador. A 50 kg sample, with a d80 of 10 cm, was initially crushed in a closed circuit using a BICO Inc. jaw crusher (Quito, Ecuador) and an ASEA HM 4100 roller crusher (Quito, Ecuador) to achieve a d80 of 2 mm. The material was then homogenized and quartered. Finally, the ore was ground in a ball mill at 62.5% solids for 40 min.
For mineralogical characterization and leaching tests, several analyses were performed. The elemental composition was determined using X-ray Fluorescence (XRF) with a Bruker S8 Tiger (Bruker, Karlsruhe, Germany). The mineralogical composition was analyzed via X-ray Diffraction (XRD) using a Bruker AXS D8 Advance (Bruker, Karlsruhe, Germany). The copper content was measured using Atomic Absorption Spectroscopy (AAS) (Perkin Elmer AA 300, Shelton, CT, USA), and precious metals (Au, Ag) were estimated through a fire assay, followed by dore dissolution for AAS measurement.
The concentration of dissolved elements (Au, Ag, Cu, Fe, and Pb) during all leaching tests was measured using AAS (Perkin Elmer AA 300).

2.2.2. Sodium Cyanide Leaching

As a reference for gold recovery, a cyanidation test was conducted using 100 g of ore with a sodium cyanide concentration of 1 (g L−1) (technical grade) for 24 h to compare the recoveries with those obtained from thiosulfate systems. The test conditions were as follows: ambient temperature (20 °C), pH 12, 750 RPM, and 15% solids. After the test, the sample was filtered using a vacuum filter, and both the pregnant solution and wash solution were analyzed to determine the gold concentration by AAS (Perkin Elmer AA 300).

2.2.3. Ammonium Thiosulfate Leaching

As a reference for gold recovery with the traditional ammoniacal thiosulfate system, tests were conducted using 100 g of ore at 15% solids for 6 h with mechanical stirring at 750 RPM. The concentration of sodium thiosulfate (sodium thiosulfate anhydrous, LOBA Chemie, 98%) was varied between 0.3 and 1 M, the pH ranged from 10 to 12, the copper sulfate concentration was adjusted between 2 and 10 mM, and the temperatures were tested between 20 and 60 °C (Figure 1a) to determine the optimal conditions for gold recovery.
At the beginning of each test, the pH was adjusted with a 20% (v v−1) ammoniacal solution, increasing the ammonium ion concentration to form the cuprotetramine complex [11,13,33,34]. Upon completion of the tests, the samples were filtered through a vacuum filter; the solution was analyzed using AAS (Perkin Elmer AA 300), and the dried ore sample was subjected to fire assay to complete the gold recovery balance.
During the tests, 10 mL aliquots were taken at 1, 2, and 4 h, which were replenished with distilled water. One part of the filtered solution was used for AAS (Perkin Elmer AA 300) to analyze the leaching kinetics, while the other part was used to determine the concentration of free thiosulfate by iodometric titration. The consumed thiosulfate was weighed and replenished to maintain the initial concentration of each test.

2.2.4. Thiosulfate–Glycine Leaching

Leaching tests with the thiosulfate–glycine system were conducted to determine the effectiveness of glycine as a replacement for ammonia. The tests were carried out in a container covered with aluminum due to the photosensitivity of glycine in the presence of potassium permanganate. The tests used 100 g of mineral at 15% solids for 6 h under mechanical agitation at 750 RPM. The variables considered included the following: sodium thiosulfate concentration from 0.3 to 0.7 M, glycine concentration (Glycine, Sigma-Aldrich, 99%) from 0.3 to 1.75 M, copper sulfate concentration at 2, 5, and 10 mM, potassium permanganate concentration at 0.004 M and 0.04 M, pH 9.3, and a temperature of 60 °C, as shown in Figure 1b, to achieve the optimal gold dissolution. At the end of the test, the sample was filtered using a vacuum filter, the solution was sent for AAS (Perkin Elmer AA 300), and the dried sample was subjected to the fire assay to quantify gold recovery.
pH adjustment was performed at the beginning of the test by adding a 4 M NaOH solution. Additionally, potassium permanganate was tested as a catalyst in the formation of the copper–glycinate complex [33,35,36] and in the co-leaching of gold due to the action of glycine [1,29,37]. The effectiveness of potassium permanganate was evaluated by increasing its concentration to 0.04 M at the start of the test and adding lower concentrations (0.004 M) at intervals of 0, 2, and 4 h.
Finally, as in the ammoniacal thiosulfate tests, 10 mL aliquots were taken at 1, 2, and 4 h and replaced with distilled water. One part of the filtered solution was used for atomic absorption to analyze the leaching kinetics, and the second part was used to determine the thiosulfate concentration through iodometric titration.

2.2.5. Glycine Leaching

A 100 g sample with 15% solids was used for 6 h of mechanical agitation at 750 RPM, with a pH of 9.3, temperature of 60 °C, and a glycine concentration of 1.75 M. Additionally, potassium permanganate was added to reach a concentration of 0.004 M at the beginning, and at 2 and 4 h.
The pH of the glycine solution, compared to the thiosulfate–glycine system tests, showed lower NaOH consumption to reach a pH of 10.5. During the test, aliquots of the solution were taken at 1, 2, and 4 h to establish the leaching kinetics and glycine concentration.
At the end of the test, the sample was filtered using a vacuum filter, the solution was sent for atomic absorption spectroscopy analysis, and the dried mineral was subjected to the fire assay.

3. Results

3.1. Characterization of the Auriferous Ore

The ore used in this study was obtained from the Guayzimi mining area in Zamora Chinchipe. The mineralogical analysis, presented in Table 1 and Figure 2 and conducted through X-ray diffraction (XRD), revealed that the primary mineral composition consists of Grosular (64%), Clinochlore (12%), and quartz (10%).
Table 2 shows the concentration of valuable metals along with the ore chemical composition, determined through X-ray fluorescence (XRF), fire assay, and acid digestion. The ore has a head grade of 10 (g t−1) of gold with a copper content of 0.008%. Other elements detected include iron (5%), calcium (15.1%), silicon (18.8%), and sulfur (0.1%). Chalcopyrite was also observed in hand sample geological assessments; however, its low concentration (>1%) prevents detection by XRD.

3.2. Sodium Cyanide Leaching

A cyanidation test was conducted to establish a comparative baseline for gold recovery. As shown in Figure 3, under the conditions of 15% solids, pH 11 (adjusted with NaOH), and a sodium cyanide concentration of 1 g (NaCN) L−1, a gold recovery of 89% was achieved after 24 h.

3.3. Ammoniacal Thiosulfate–Copper Leaching

Leaching tests using the ammoniacal thiosulfate–copper system were conducted to evaluate gold recovery under different leaching conditions. The optimized conditions were as follows: sodium thiosulfate concentration of 0.7 M, pH 10.5 (adjusted with a 20% v v−1 ammoniacal solution), copper sulfate concentration of 5 mM, 750 RPM agitation, and 15% solids for 6 h. Under these conditions, a maximum gold recovery of 40% was achieved, as shown in Figure 3.
Additionally, a reduction in thiosulfate concentration to 0.5 M was observed during the first 2 h of testing. Consequently, further tests were conducted by systematically varying the conditions, including thiosulfate concentration (0.5–1 M), copper sulfate concentration (2–10 mM), and pH (10.5–11.5) to optimize gold recovery from the characterized ore using the ammoniacal thiosulfate–copper system.

Evaluation of Leaching Conditions for Gold Recovery Using the Ammoniacal Thiosulfate System

Leaching tests were performed by varying cupric ion concentration, pH, and temperature to assess the efficiency of the ammoniacal thiosulfate system. Results indicate that reducing the copper sulfate concentration to 2 mM increases gold recovery to 48%, as shown in Figure 4a, whereas a higher concentration of 10 mM only achieves a gold dissolution rate of 33%.
Additionally, increasing the ammonia content to raise the pH from 10.5 to 11.5 resulted in an improvement in gold recovery from 40% to 50%, as depicted in Figure 4b.
Finally, tests conducted at different temperatures revealed that raising the temperature to 60 °C enhances recovery to 58% in 6 h, as presented in Figure 5.
The optimal leaching conditions for the ammoniacal thiosulfate system were established at 0.7 M sodium thiosulfate, 2 mM copper sulfate, pH 10.5, and a temperature of 60 °C for 6 h, achieving a gold recovery of 58%.

3.4. Thiosulfate–Glycine–Copper Leaching

In the baseline test with the thiosulfate–glycine–copper system, using 100 g of mineral with a 15% solid concentration, a gold recovery of 40% was achieved under the following conditions: 0.7 M sodium thiosulfate, 0.3 M glycine, 60 °C, agitation at 750 RPM, and pH 10.5 (adjusted with NaOH) over 6 h, as shown in Figure 6a. This test served as a reference point for subsequent experimental variations, including adjustments in sodium thiosulfate concentration, glycine concentration, copper sulfate pentahydrate, potassium permanganate, and pH, aiming to enhance gold recovery efficiency.

3.4.1. Effect of Sodium Thiosulfate Concentration

Table 3 presents the results of tests in which the sodium thiosulfate concentration was varied (0.5 M, 0.7 M, and 1 M), while all other experimental conditions remained constant. As shown in Figure 6a, gold recovery increased proportionally with the sodium thiosulfate concentration. The highest recovery, 49%, was achieved with a concentration of 1 M.
However, the difference in gold recovery between 1 M (49%) and 0.7 M (40.8%) was not substantial enough to justify the increase in reagent concentration. Thus, a 0.7 M concentration was selected for subsequent tests.
Throughout the tests, iodometric titration of the thiosulfate concentration showed no consumption of reagent over the 6 h of agitation.

3.4.2. Effect of Copper Ion Concentration

Figure 6b illustrates gold recovery in the absence of added copper sulfate, indicating that dissolved copper from the mineral in the presence of glycine was insufficient, yielding a recovery of less than 30%. To improve this outcome, tests were conducted with varying copper sulfate concentrations of 2 mM, 5 mM, and 10 mM, as detailed in Figure 6b.
Increasing the copper sulfate concentration to 5 mM enhanced gold dissolution, achieving a recovery of 48%. However, raising the concentration to 10 mM did not further improve recovery efficiency. The excess of cupric ions accelerated thiosulfate degradation, as evidenced by a reduction in thiosulfate concentration to 0.65 M when using 10 mM copper sulfate during the first 2 h of testing, quantified through iodometry.
In the thiosulfate–glycine system, no thiosulfate consumption was detected throughout the test when copper sulfate concentrations were kept below 5 mM.

3.4.3. Effect of Glycine Concentration

Leaching tests were conducted with a sodium thiosulfate concentration of 0.7 M and copper sulfate at 5 mM, varying the glycine concentration between 0.3 M, 0.5 M, 1 M, and 1.75 M. The results, shown in Figure 7, demonstrate that an increase in glycine concentration enhances gold recovery, achieving a dissolution of 69% of the gold with a glycine concentration of 1.75 M after 6 h.

3.4.4. Effect of pH

Although previous studies [5,13] suggest that optimal gold recovery in a thiosulfate system is achieved at pH levels above 10.5, the results in Figure 8 indicate that gold recovery is equivalent at pH 9.3 and 10.5, reaching 65% in both cases. These tests were conducted with the same reagent concentrations. It was observed that glycine concentration is directly related to NaOH consumption, requiring larger volumes of NaOH to maintain a pH of 10.5 when the glycine concentration exceeds 1 M.

3.4.5. Effect of Oxidizing Agents: Potassium Permanganate

Tests were conducted to evaluate the effect of potassium permanganate as an oxidizing agent in the gold-leaching process using the thiosulfate–glycine system. As shown in Figure 9a, the addition of potassium permanganate at a concentration of 0.004 M improved the gold recovery, achieving 65% recovery after 6 h.

Effect of Potassium Permanganate Concentration

Tests with an initial potassium permanganate concentration of 0.004 M resulted in a 65% gold recovery, as shown in Figure 9a. However, when the concentration was increased to 0.04 M, gold recovery decreased to 55%. Additionally, thiosulfate consumption was observed during the first 2 h.

Effect of Periodic Addition of Potassium Permanganate

Figure 9b shows the results of tests in which potassium permanganate at 0.004 M was periodically added at the start, at 2 h, and at 4 h. This approach achieved a 70% gold recovery after 6 h, maintaining a constant sodium thiosulfate concentration throughout the test.
Additionally, to assess the efficiency of glycine alone in dissolving gold, tests were conducted using only glycine at a concentration of 1.75 M, which resulted in a 6% gold dissolution. When potassium permanganate was added to the glycine, the gold recovery increased to 22%, as illustrated in Figure 10.

3.4.6. Synergy of Conditions for Gold Recovery with a Thiosulfate–Glycine–Copper-Permanganate System

Tests conducted under optimal conditions 0.7 M sodium thiosulfate, 1.75 M glycine, 5 mM copper sulfate, 0.004 M potassium permanganate added periodically, pH 9.3, 60 °C, and 750 RPM achieved an 89% gold recovery after 6 h, as shown in Figure 11. In the initial four control tests, gold recovery reached 85% within the first 4 h of leaching.

4. Discussion

4.1. Characterization of Auriferous Ore

The mineralogical characterization revealed that the ore is composed of 78% aluminosilicates (Table 1), suggesting that the gold is trapped within this matrix, particularly associated with clinochlore and grossular.
The copper content, at 0.008% (Table 2), acts as a catalyst in the dissolution of gold in ammoniacal-thiosulfate-leaching systems [3,5]. These findings highlight the importance of leaching systems that leverage the interaction between copper and thiosulfate to improve gold recovery. Furthermore, the presence of chalcopyrite, albeit at low concentration, can influence the kinetics of gold leaching, interferes with the dissolution process, and reduces the ability to form the thiosulfate gold complex.

4.2. Gold Recovery by Cyanidation Leaching

Cyanidation, although effective, achieved an 89% recovery for this mineral, which differs from the 98% recoveries reported in previous studies [1,2]. This recovery is explained by the mineralogical association of gold with aluminosilicates, specifically clinochlore and grossular, which limits the complete release of the metal. This result will be used as a reference to compare the recovery achieved with other leaching systems.
Additionally, the formation of copper cyanide complexes competes with gold, decreasing the concentration of dissolved sodium cyanide [1,2]. These results demonstrate the need to consider alternative methods that utilize copper as a catalytic agent, such as sodium thiosulfate.

4.3. Ammoniacal Thiosulfate-Copper Leaching System

Leaching with ammoniacal thiosulfate and copper resulted in only a 40% gold recovery, compared to the expected 85% under optimal conditions [13]. This low efficiency can be explained by the rapid degradation of thiosulfate into products such as polythionates, reducing its availability to form a stable complex with gold. This phenomenon was observed within the first 2 h of testing, as noted in other studies [38,39].
Furthermore, factors such as cupric ion concentration above 5 mM, pH, temperature, and the mineralogical composition of the ore accelerate the degradation of thiosulfate. The copper in the ore acts as a catalyst in forming the cuprotetramine complex, which impacts thiosulfate stability. These results suggest that, while the ammoniacal thiosulfate–copper system is a viable alternative, optimizing operational conditions is essential to minimize thiosulfate degradation and improve gold recovery.

Evaluation of Leaching Conditions for Gold Recovery Using the Ammoniacal Thiosulfate System

Tests with ammoniacal thiosulfate demonstrated that a lower copper sulfate concentration (2 mM) increases gold recovery from 38% to 48%. In contrast, higher copper concentrations (>5 mM) reduce gold recovery to 33%. This is because concentrations exceeding 5 mM accelerate thiosulfate degradation, limiting its effectiveness as a leaching agent. These findings align with reports by Aylmore and Muir (2001) [38], who observed reduced gold recovery with excessive copper concentrations in the leaching system.
Moreover, increasing the solution pH due to higher ammonia concentration favors the formation of the cupric–tetramine complex, stabilizing cupric ions, as noted by Senanayake (2007) [40] and Xu et al. (2017) [41]. This effect was observed when increasing the pH from 10.5 to 11.5 (Figure 12), where recovery increased from 40% to 50%. However, it is important to highlight that higher ammonia concentrations present environmental risks [42].
Finally, tests conducted at 60 °C and 2 mM of copper sulfate showed a maximum recovery of 58% (Figure 12), confirming that temperature has a direct effect on enhancing gold dissolution in the ammoniacal thiosulfate system, consistent with previous studies by Senanayake (2005) [31] and Aylmore and Muir (2001) [38]. These results suggest that temperature (60 °C) is a critical factor for optimizing the leaching process and achieving higher gold recovery.

4.4. Thiosulfate–Glycine–Copper-Leaching System

The baseline test using the thiosulfate–glycine–copper system achieved an initial gold recovery of 38%, which is promising to consider that it was conducted at 60 °C and with a low glycine concentration (0.3 M). Glycine, as mentioned in previous research [24], has the potential to form a stable cupric complex, facilitating gold leaching. However, process efficiency needs improvement.
Variations in the concentration of thiosulfate, glycine, copper sulfate, potassium permanganate, and pH are expected to accelerate gold dissolution (Figure 13). Higher concentrations of these reagents may improve the stability of cupric ions by forming the copper glycinate complex, preventing thiosulfate degradation and increasing gold dissolution, as noted by Barani et al. (2022) [11] and Oraby and Eksteen (2015) [29]. Moreover, the addition of potassium permanganate as an oxidizing agent could accelerate gold leaching in this system.
The temperature for all experiments was maintained at 60 °C, based on the results obtained with the ammoniacal system at this temperature and to prevent glycine degradation at temperatures above 60 °C [43].

4.4.1. Effect of Sodium Thiosulfate Concentration

Increasing the sodium thiosulfate concentration results in improved gold recovery, as expected, since a higher thiosulfate concentration facilitates the formation of the gold–thiosulfate complex, in line with the studies by Gámez et al. (2019) [7]. However, the difference in gold recovery between 0.7 M (41%) and 1 M (49%) is not significant enough to justify the use of a higher reagent concentration.
Additionally, the absence of thiosulfate consumption during tests indicates that glycine forms more stable cupric complexes than the cuprotetramine complex. This is consistent with stability studies presented by Wang (2022) [12].

4.4.2. Effect of Copper Ion Concentration

The concentration of dissolved cupric ions is a key factor in gold recovery in thiosulfate systems. The results show that a copper sulfate concentration of 5 mM is optimal for promoting gold dissolution, increasing efficiency without compromising thiosulfate stability. However, an excess copper sulfate concentration (10 mM) accelerated thiosulfate degradation, reducing its concentration to 0.65 M during the first two hours of the test.
In comparison to the ammoniacal thiosulfate system, which showed significant thiosulfate consumption even at low copper concentrations, the thiosulfate–glycine system demonstrates greater stability. This is attributed to glycine’s ability to form more stable complexes, such as the copper glycinate, as also reported in previous studies [1,29,44].

4.4.3. Effect of Glycine Concentration

A gold recovery of 69% obtained with a glycine concentration of 1.75 M demonstrates that glycine plays a key role in improving the gold recovery process. The observed trend, where gold recovery increases proportionally with glycine concentration, can be explained by its dual capability: forming the copper glycinate complex, which catalyzes the formation of the gold–thiosulfate complex, and acting directly as a co-leaching agent, contributing to gold dissolution.
These results are consistent with previous studies suggesting that glycine, by stabilizing cupric ions through the formation of the copper glycinate complex, enhances thiosulfate stability and additionally facilitates gold lixiviation, thereby improving the overall efficiency of the gold recovery process [1,22,29].

4.4.4. Effect of pH on Thiosulfate–Glycine System

Although previous studies indicate that a pH close to 10.5 optimizes gold recovery in sodium-thiosulfate-leaching systems [45,46], this study demonstrated that maintaining a pH of 9.3 is sufficient to achieve similar results (Figure 8). This is because glycine has the ability to stabilize cupric ions over a broader pH range, reducing the need to operate under highly alkaline conditions.
The formation of a ‘buffer’ at pH 9.4, attributed to the properties of glycine [47,48,49], explains the increased NaOH consumption observed at glycine concentrations above 1 M when adjusting the pH to values above 10. Nevertheless, for the thiosulfate–glycine–copper system, maintaining a pH of 9.3 is sufficient to achieve 65% gold recovery.

4.4.5. Effect of Oxidizing Agents: Potassium Permanganate

The presence of potassium permanganate accelerates gold dissolution in the thiosulfate–glycine system, consistent with findings by Eksteen and Oraby (2015) [24] and Senanayake (2012) [45], who reported significant improvements in gold recovery when using additional oxidants. Potassium permanganate’s oxidizing effect creates a more favorable environment for glycine’s role as a co-leaching agent, promoting the formation of stable complexes with copper and gold and enhancing the gold recovery.

Effect of Potassium Permanganate Concentration

The results indicate that increasing the potassium permanganate concentration does not improve gold recovery. In fact, increasing the concentration from 0.004 M to 0.04 M resulted in a decrease in final gold recovery from 65% to 55%. This demonstrates that an excessively oxidizing environment accelerates thiosulfate degradation.
This phenomenon aligns with previous studies indicating that excessive oxidant concentrations can break down thiosulfate into unwanted products, reducing the availability of thiosulfate [24,45]. Thiosulfate consumption observed during the first 2 h of testing with a concentration of 0.04 M confirmed this.

Effect of Periodic Addition of Potassium Permanganate

The results confirmed that periodic addition of potassium permanganate improves gold recovery in the thiosulfate–glycine system. The improvement in gold recovery to 70%, compared to the 65% obtained with a single addition of permanganate at the start of the test, suggests that the effective action time of permanganate as an oxidizing agent is approximately 2 h. This aligns with previous research indicating that a continuous and controlled oxidant concentration prevents premature thiosulfate degradation [38,45,50].
Although glycine with permanganate increases the gold dissolution rate, Figure 10 shows that glycine alone is inefficient (6% recovery), suggesting that its role is more effective in stabilizing copper complexes within the thiosulfate system rather than acting as an independent leaching agent. Even with potassium permanganate, glycine alone does not achieve acceptable recovery (22%), reinforcing that glycine’s primary function is to stabilize cupric ions, accelerating gold dissolution, rather than acting as an independent lixiviant.

4.4.6. Synergy of Conditions for Gold Recovery with a Thiosulfate–Glycine–Copper–Permanganate System

The results demonstrate that the synergy between thiosulfate, glycine, copper, and permanganate improves gold dissolution, achieving an 89% recovery in 6 h. The efficiency observed in the tests suggests that this system can recover most of the gold within the first 4 h, these results are consistent with previous studies [27] that propose the combination of thiosulfate and glycine to improve leaching kinetics.
Using glycine to stabilize cupric ions, along with periodic addition of potassium permanganate, is key to accelerating the process without significant thiosulfate degradation. Additionally, the results indicate that reducing leaching time does not affect system efficiency, suggesting that this combination of reagents is both effective and economical in terms of time and reagent consumption [27].
If the leaching system is to be scaled up to an industrial level, the primary challenges include maintaining precise temperature control and protecting the solution from sunlight during the process. Tests with permanganate revealed the photosensitivity of the solution, leading to the generation of byproducts that significantly reduced gold recovery. To mitigate this issue, aluminum covers were required to shield the containers during agitation.

4.4.7. Mechanism of Gold Dissolution in the Thiosulfate–Glycine–Copper–Permanganate System

As shown in Figure 14, the mechanism of gold dissolution in the thiosulfate–glycine–copper–permanganate system relies on their synergistic interaction in a moderately alkaline environment (pH 9.3) and at a temperature of 60 °C. The general reaction involved in gold dissolution is presented in Equation (1). Based on experimental assays and analyzed thermodynamic diagrams [27], for the dissolution mechanism, the presence of two electrochemical areas between the gold surface and the leaching solution was assumed: the cathodic area and the anodic area.
Au + 2(S2O3)2− + KMnO4 + Cu(C2H2NO2)3+H2O → Au(S2O3)23− + Cu(C2H2NO2)3 + MnO2 + O2 + OH−
In the cathodic area, glycine plays a dual role. First, it forms the cupric glycinate complex ([Cu(C2H4NO2)2]0), which remains stable due to the high affinity between glycine and copper [27]. Through a REDOX reaction, this cupric glycinate complex interacts with dissolved thiosulfate ions to form the Cu(S2O3)35− complex via electron transfer, reducing copper (II) to copper (I). The continuous regeneration of the cupric glycinate complex is facilitated by the oxidizing environment created by dissolved oxygen and potassium permanganate. This ensures the constant oxidation of Cu(I) back to Cu(II) [27]. This process not only stabilizes the cupric ions in the glycinate complex but also prevents the accumulation of unstable species that could degrade thiosulfate into polythionates. Additionally, free glycine produced during the REDOX regeneration process can form a gold–glycinate complex.
In the anodic area, cupric ions stabilized as cupric glycinate facilitate the direct action of thiosulfate on metallic gold, resulting in the formation of the gold–thiosulfate complex Au(S2O3)23−. The stability of the cupric glycinate in the medium prevents thiosulfate decomposition, as confirmed experimentally by the absence of reagent consumption. Furthermore, although the formation of a gold–glycinate complex is possible, this complex is thermodynamically less stable than the gold–thiosulfate complex [22,27]. Therefore, in the presence of thiosulfate, gold initially associated with glycine transitions to the gold–thiosulfate complex Au(S2O3)23−, regenerating glycine for further complexation in the cathodic area during the REDOX cycle.
Finally, experimental assays demonstrated that in the presence of glycine and permanganate, no thiosulfate consumption was recorded. This supports the conclusion that in an oxidizing environment, glycine’s primary role is to stabilize copper in its Cu(II) state (Figure 14). This stabilization prevents the formation of unstable cuprous complexes that accelerate thiosulfate degradation. Although glycine has the capacity to form complexes with gold, experimental results show that its primary interaction is with copper. Glycine stabilizes copper in its cupric state, maintaining an optimal environment for thiosulfate to dissolve gold (Figure 14). This reaffirms that glycine’s primary role is associated with copper rather than gold.

5. Conclusions

The auriferous ore from the mining area in Ecuador exhibited a head grade of 10 (g t−1) of gold and 0.008% Cu. The main mineralogy obtained through the XRD analysis revealed aluminosilicates, such as grossular (64%) and clinochlore (12%), to which the gold is associated. Additionally, the chemical analysis performed via XRF indicated a composition of 18.8% Si, 15.1% Ca, 5% Fe, and 4.5% Al, with trace amounts of manganese (1%), sulfur (0.1%), and titanium (0.2%).
Baseline cyanidation tests achieved a gold recovery of 89.8% at room temperature in 24 h, highlighting the process’s efficiency. In contrast, ammoniacal thiosulfate leaching reached a maximum recovery of 53.1% under optimal conditions (0.7 M thiosulfate, 2 mM copper sulfate, pH 10.5, and 60 °C). However, the degradation of thiosulfate accelerated with copper sulfate concentrations above 2 mM due to the increased availability of cupric ions generated by copper dissolution from the ore in the presence of ammonia.
The thiosulfate–glycine system exhibited an initial recovery of 40%, and it was determined that a sodium thiosulfate concentration of 0.7 M is sufficient, as higher concentrations do not significantly improve recovery. Regarding copper requirements, the glycine system showed greater stability compared to the ammoniacal system, achieving a gold recovery of 47% with 5 mM copper sulfate and no significant consumption of sodium thiosulfate. This can be attributed to the ability of glycine to form more stable cupric complexes compared to cupric–tetramine, although its capacity to dissolve copper from the ore is lower than that of ammonia. Copper concentrations above 10 mM were counterproductive, yielding the same recoveries as at 5 mM and accelerating the degradation and consumption of thiosulfate during the first four hours of agitation.
The glycine concentration also proved crucial for stabilizing dissolved cupric ions, preventing thiosulfate degradation, and acting as a co-leaching agent for gold. A glycine concentration of 1.75 M resulted in the highest recovery of 60%. Additionally, the system maintained stable recoveries of approximately 65% at both pH 10.5 and pH 9.3, due to glycine’s ability to form stable cupric complexes over a broader pH range compared to ammonia. This behavior minimizes the need to adjust the pH to values above 10, given the buffering effect observed at pH 9.4.
The use of potassium permanganate as an oxidizing agent at a concentration of 0.004 M accelerated the co-leaching action of glycine (1 M), improving gold recovery to 65% in 6 h. However, higher concentrations (0.04 M) decreased recovery to 55%, as an excessively oxidizing environment, while initially enhancing gold dissolution, also accelerated thiosulfate degradation, resulting in reagent consumption after the second hour of agitation. The periodic addition of potassium permanganate (0.004 M) every two hours maintained a lower and more stable oxidizing environment, enabling a gold recovery of 70% in 6 h. In contrast, tests conducted solely with glycine (1.75 M) and periodically added permanganate (0.004 M) achieved a limited recovery of 22%, underscoring the importance of the synergy between thiosulfate, glycine, copper, and permanganate.
Finally, the optimized thiosulfate–glycine–copper–permanganate system achieved a gold recovery of 89.3% under the following conditions: 0.7 M sodium thiosulfate, 1.75 M glycine, 5 mM copper sulfate, 0.004 M potassium permanganate added periodically every 2 h, pH 9.3, and 60 °C. Although the total leaching time was 6 h, 85% of the gold was recovered within the first 4 h, demonstrating the efficiency of the system. This performance is attributed to glycine’s ability to stabilize cupric ions, preventing thiosulfate degradation, and its role as a co-leaching agent in an oxidizing environment controlled by potassium permanganate.

Author Contributions

Conceptualization, E.d.l.T.; methodology, E.d.l.T. and A.S.R.; validation, A.S.R. and E.d.l.T.; formal analysis, A.S.R.; investigation, A.S.R.; resources, A.S.R.; data curation, A.S.R.; writing—original draft preparation, A.S.R. and C.F.A.-T.; writing—review and editing, A.S.R. and C.F.A.-T.; visualization, A.S.R.; supervision, E.d.l.T.; project administration, E.d.l.T.; funding acquisition, E.d.l.T. All authors have read and agreed to the published version of the manuscript.

Funding

The research presented in this study was made possible by the financing of the Department of Extractive Metallurgy (DEMEX) of the Escuela Politécnica Nacional thanks to the research project PIGR-2208. Tel.: +593-2-297-6300 (ext. 5806).

Data Availability Statement

The original contributions presented in this study are included in the article. Further inquiries can be directed to the corresponding author.

Conflicts of Interest

The authors declare no conflicts of interest.

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Figure 1. Comparison between the methodology followed for the (a) ammoniacal thiosulfate and (b) thiosulfate–glycine system.
Figure 1. Comparison between the methodology followed for the (a) ammoniacal thiosulfate and (b) thiosulfate–glycine system.
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Figure 2. XRD pattern for the characterization of the auriferous ore from Guayzimi-Ecuador, identifying grossular (orange), clinochlore (blue), quartz (green), and calcite (purple) as the principal minerals. Other minerals are not highlighted as their concentrations do not exceed 5%.
Figure 2. XRD pattern for the characterization of the auriferous ore from Guayzimi-Ecuador, identifying grossular (orange), clinochlore (blue), quartz (green), and calcite (purple) as the principal minerals. Other minerals are not highlighted as their concentrations do not exceed 5%.
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Figure 3. Comparison between baseline cyanidation tests (1 g L−1; 15% solids; pH 12 for 24 h) and ammoniacal thiosulfate tests (Na2S2O3 0.7 M; copper sulfate 5 mM; pH 10.5; 15% solids for 6 h).
Figure 3. Comparison between baseline cyanidation tests (1 g L−1; 15% solids; pH 12 for 24 h) and ammoniacal thiosulfate tests (Na2S2O3 0.7 M; copper sulfate 5 mM; pH 10.5; 15% solids for 6 h).
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Figure 4. (a) Effect of copper concentration on Au recovery using the ammoniacal thiosulfate system: Na2S2O3 (0.7 M); pH (10.5). (b) Effect of pH on gold recovery using the ammoniacal thiosulfate system: Na2S2O3 (0.7 M); copper sulfate (2 mM).
Figure 4. (a) Effect of copper concentration on Au recovery using the ammoniacal thiosulfate system: Na2S2O3 (0.7 M); pH (10.5). (b) Effect of pH on gold recovery using the ammoniacal thiosulfate system: Na2S2O3 (0.7 M); copper sulfate (2 mM).
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Figure 5. Effect of temperature on gold recovery using the ammoniacal thiosulfate system: Na2S2O3 (0.7 M); copper sulfate (2 mM) in 6 h.
Figure 5. Effect of temperature on gold recovery using the ammoniacal thiosulfate system: Na2S2O3 (0.7 M); copper sulfate (2 mM) in 6 h.
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Figure 6. (a) Effect of sodium thiosulfate concentration on gold recovery using the thiosulfate–glycine system: C2H5NO2 (0.3 M); 60 °C; pH (10.5); copper sulfate (2 mM). (b) Effect of copper concentration on gold recovery using the thiosulfate–glycine system: Na2S2O3 (0.7 M); glycine (0.3 M); 60 °C; pH (10.5).
Figure 6. (a) Effect of sodium thiosulfate concentration on gold recovery using the thiosulfate–glycine system: C2H5NO2 (0.3 M); 60 °C; pH (10.5); copper sulfate (2 mM). (b) Effect of copper concentration on gold recovery using the thiosulfate–glycine system: Na2S2O3 (0.7 M); glycine (0.3 M); 60 °C; pH (10.5).
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Figure 7. Effect of glycine (C2H5NO2) concentration on Au recovery using the thiosulfate–glycine system: Na2S2O3 (0.7 M); copper sulfate (5 mM); 60 °C; pH (10.5); 15% solids for 6 h.
Figure 7. Effect of glycine (C2H5NO2) concentration on Au recovery using the thiosulfate–glycine system: Na2S2O3 (0.7 M); copper sulfate (5 mM); 60 °C; pH (10.5); 15% solids for 6 h.
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Figure 8. Effect of pH on gold recovery using the thiosulfate–glycine system: Na2S2O3 (0.7 M); C2H5NO2 (1 M); copper sulfate (5 mM); 60 °C; 15% solids for 6 h.
Figure 8. Effect of pH on gold recovery using the thiosulfate–glycine system: Na2S2O3 (0.7 M); C2H5NO2 (1 M); copper sulfate (5 mM); 60 °C; 15% solids for 6 h.
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Figure 9. (a) Effect of potassium permanganate concentration on gold recovery using the thiosulfate–glycine system: Na2S2O3 (0.7 M); C2H5NO2 (1 M); copper sulfate (5 mM); 60 °C; pH (9.3); 15% solids for 6 h. (b) Effect of periodic addition of potassium permanganate on gold recovery using the thiosulfate–glycine system: Na2S2O3 (0.7 M); C2H5NO2 (1 M); copper sulfate (5 mM); 60 °C; pH (9.3); 15% solids for 6 h.
Figure 9. (a) Effect of potassium permanganate concentration on gold recovery using the thiosulfate–glycine system: Na2S2O3 (0.7 M); C2H5NO2 (1 M); copper sulfate (5 mM); 60 °C; pH (9.3); 15% solids for 6 h. (b) Effect of periodic addition of potassium permanganate on gold recovery using the thiosulfate–glycine system: Na2S2O3 (0.7 M); C2H5NO2 (1 M); copper sulfate (5 mM); 60 °C; pH (9.3); 15% solids for 6 h.
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Figure 10. Gold recovery using glycine and potassium permanganate solutions (PP): C2H5NO2 (1.75 M); potassium permanganate (PP) (0.004 M); 60 °C; pH (9.3); 15% solids for 6 h.
Figure 10. Gold recovery using glycine and potassium permanganate solutions (PP): C2H5NO2 (1.75 M); potassium permanganate (PP) (0.004 M); 60 °C; pH (9.3); 15% solids for 6 h.
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Figure 11. Gold recovery using the thiosulfate–glycine system: Na2S2O3 (0.7 M); C2H5NO2 (1.75 M); copper sulfate (5 mM); potassium permanganate 0.004 M added every two hours; 60 °C; pH (9.3); 15% solids for 6 h.
Figure 11. Gold recovery using the thiosulfate–glycine system: Na2S2O3 (0.7 M); C2H5NO2 (1.75 M); copper sulfate (5 mM); potassium permanganate 0.004 M added every two hours; 60 °C; pH (9.3); 15% solids for 6 h.
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Figure 12. Stepwise optimization of pH, copper sulfate concentration, and temperature for improved gold recovery (40% to 58%).
Figure 12. Stepwise optimization of pH, copper sulfate concentration, and temperature for improved gold recovery (40% to 58%).
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Figure 13. Sequential optimization of thiosulfate, copper sulfate, glycine, pH, potassium permanganate concentration, and dosing frequency to achieve 89% gold recovery.
Figure 13. Sequential optimization of thiosulfate, copper sulfate, glycine, pH, potassium permanganate concentration, and dosing frequency to achieve 89% gold recovery.
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Figure 14. Proposed electrochemical mechanism of gold dissolution in the thiosulfate–glycine–copper–permanganate system. The central arrow represents the electron exchange between the anodic and cathodic areas, while the highlighted boxes indicate the key chemical species, cupric glycinate and gold thiosulfate, targeted for formation within the system.
Figure 14. Proposed electrochemical mechanism of gold dissolution in the thiosulfate–glycine–copper–permanganate system. The central arrow represents the electron exchange between the anodic and cathodic areas, while the highlighted boxes indicate the key chemical species, cupric glycinate and gold thiosulfate, targeted for formation within the system.
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Table 1. Mineralogical characterization of Guayzimi ore by XRD analysis.
Table 1. Mineralogical characterization of Guayzimi ore by XRD analysis.
MineralFormulaContent (%)
GrossularCa3Al2(SiO4)364
Clinochlore(Mg, Fe)5Al (Si, Al)4O10 (OH)812
QuartzSiO210
CalciteCaCO38
KaoliniteAl2(Si2O5) (OH)44
Plagioclases(Na, Ca) Al (Si, Al) Si2O82
Table 2. Elemental composition, gold, and copper content of the ore determined using XRF, fire assay, and acid digestion.
Table 2. Elemental composition, gold, and copper content of the ore determined using XRF, fire assay, and acid digestion.
ElementConcentration
Au10 g t−1
Cu0.008%
Si18.8%
Ca15.1%
Fe5%
Al4.5%
Mg1.6%
Mn0.9%
K0.8%
Na0.4%
S0.1%
Table 3. Effect of sodium thiosulfate concentration on gold recovery using the thiosulfate–glycine system after 6 h. Glycine 0.3 M, 60 °C, pH 10.5.
Table 3. Effect of sodium thiosulfate concentration on gold recovery using the thiosulfate–glycine system after 6 h. Glycine 0.3 M, 60 °C, pH 10.5.
Sodium Thiosulfate ConcentrationGold Recovery (%)
0.5 M33.7
0.7 M40.8
1 M49.6
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Redrovan, A.S.; Torre, E.d.l.; Aragón-Tobar, C.F. Gold Leaching from an Auriferous Ore by Alkaline Thiosulfate–Glycine–Copper Solution. Metals 2025, 15, 204. https://doi.org/10.3390/met15020204

AMA Style

Redrovan AS, Torre Edl, Aragón-Tobar CF. Gold Leaching from an Auriferous Ore by Alkaline Thiosulfate–Glycine–Copper Solution. Metals. 2025; 15(2):204. https://doi.org/10.3390/met15020204

Chicago/Turabian Style

Redrovan, Alex S., Ernesto de la Torre, and Carlos F. Aragón-Tobar. 2025. "Gold Leaching from an Auriferous Ore by Alkaline Thiosulfate–Glycine–Copper Solution" Metals 15, no. 2: 204. https://doi.org/10.3390/met15020204

APA Style

Redrovan, A. S., Torre, E. d. l., & Aragón-Tobar, C. F. (2025). Gold Leaching from an Auriferous Ore by Alkaline Thiosulfate–Glycine–Copper Solution. Metals, 15(2), 204. https://doi.org/10.3390/met15020204

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